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By Jaracz K

At the operation of mine winders in deep shafts considerable oscillations of winding ropes may occur due to elasticity of both wind-ing and balance ropes. The amplitudes of longitudinal oscillations may reach + lm at relatively low frequencies, ranging from 0.5 to 2 c/s. These oscillations may be dangereus for the rope and produce several hazards to the winding installation as a whole. They may cause conside-rable instant accelerations on the pulley and dangerous tensions in ro-pe wires. The excessive acceleration may cause the uncontrolled rope slip on the pulley. Several methods of limitation of these accelera-tions were developed, some of them have been already in control system: of electric drive of the winder. The most simple ones are: a/ limita-tion of the rate of change of acceleration and retardation of the win-der /so called "jerk"/ and b/ the proper selection of regulators of the winder control system to get the optimal control. The trouble in solv-ing the problem of rope oscillations is that the dynamic system of the winder is a nonstationaryone owing to changing distance of conveyance from the pulley. Besides, the system is described by the equations with distributed parameters. The brief discussion of the theory and its practical aspects are enclosed as well.

Jan 1, 1985

Mr D. Twigg graduated in Metallurgy at the Sheffield Polytechnic, and later obtained a Master of Science degree at the University of Bradford. Mr I.D. Matthews matriculated at Roosevelt High School and joined the No. 1 Parachute Battalion in 1968. He worked in Zimbabwe for 31 years in geological prospecting, civil engineering, and open-pit mining.

Jan 1, 1985

By L. M. Turner

Perlite is a unique mineral that upon heating expands to about twenty times its original volume. Mining is by open pit method. The processing cycle involves crushing, screening and sizing the ore. An on-site quality control facility at the Manville Products No Agua mine exerts a high degree of control over every phase of the operation from blasting to finished product.

Jan 1, 1985

By R. A. Seitz

This paper presents a framework for analyzing the theory and industrial practice of coal flotation and gives some examples of its use for understanding flotation phenomena and optimizing/controlling flotation circuit performance. It is based on extensive laboratory and in plant test work conducted over the past several years by researchers at MTU as well as an analysis of the literature. Previous researchers have usually considered either the surface chemistry or the kinetics of coal flotation behavior, only rarely have these factors been considered together. However, these factors must be considered together in order to gain a clear understanding of the theory underlying coal flotation such that its practical application can be optimized. Fortunately, this complex objective can be achieved by considering the recovery and rate of recovery of the different types of particles present in the feed to coal flotation circuits and the fundamental mechanisms responsible for these recovery phenomena. The influence of three major groups of variables must be considered: control variables (reagent additions, aeration rate, etc.); disturbance variables (feed pH, percent solids, etc.); and circuit design variables (number of cells, circuit arrangement, etc.). Each of these groups of variables affects process kinetics in a clearly understandable manner, which, once determined, is extremely useful in circuit design and optimization. This paper includes a discussion of the use of the proposed framework to explain the effects of some coal characteristics on flotation performance. In addition, an example of using this analysis to optimize the performance of a coal flotation circuit using frother and collector addition levels is included.

Jan 1, 1985

By James W. Clark, William C. Dixon, Kevin L. Wiley, Thomas R. Couzens

This paper is a description of certain geologic aspects of the Inspiration mining operation. It is not a geologic description of the Inspiration deposit nor an exposition of all the geologic work. Involved in the development and exploitation of the deposit. A few specific examples of the application of geology and geostatistics to the mining process are given to illustrate the progress that has been made in making the Inspiration Mine more efficient and cost effective, and a statement of the philosophy of geologic applications is offered.

Jan 1, 1985

By J. Keith Haddock

"IntroductionThe Bienfait Mine is operated by Bienfait Coal Company Ltd., one of the Luscar Limited group of companies. Situated 5 km southeast of the town of Bienfait, Saskatchewan, it can produce 2.0 million tonnes of lignite annually in addition to 158,000 tonnes of char from a carbonization plant located on the property. Geology The economically mineable coal seams of the Estevan area are found in the Ravenscrag Formation. The lignite-bearing sands, silts and clays were deposited in Early Tertiary age, in an inland sea centred in the Williston Basin of North Dakota. The top of the Ravenscrag is an erosional surface, overlain by Tertiary or Quarternary sands and gravels or glacial till. The regional structure is a gently plunging syncline which dips to the southeast at about 0.3 degrees. There are minor undulations in the Bienfait area which are due to differential subsidence during deposition. Sporadic faulting is probably a result of glacial pressures, and preglacial and glacial channels are responsible for washout zones and the abrupt irregular subcrop found in the area. There are two economic coal seams in the Bienfait area, an uppe; seam locally known as the Roche Percee, and a lower seam, the Estevan. Both seams have been extensively mined since lignite mining began in the area in the late 1800s. The Roche Percee remains only in a small portion of the Bienfait Mine and comprises a very minor portion of the reserves. It averages 1.74 m thick under 8.1 m of overburden. The till rests on top of the Roche Percee and due to weathering the seam is generally of poorer quality. The Estevan seam comprises the majority of reserves at Bienfait. The main seam which is being worked on consists of an upper bench averaging 2.75 m, a carbonaceous parting of 0.21 m, and a lower bench of 0.92 m. Overburden depth above this seam varies from 21.0 m to 38.0 m."

Jan 1, 1985

By B. L. Seegmiller

Horizontally stratified lithologies of certain open pit or strip mines create unique stability problems for mine operations. Such problems may result when specific lithologies, such as sandstone and shale, exist in final or long-term slopes. Slope instabilities may occur when a sandstone unit directly overlies a shale unit and ground water drains into the mine along the lithologic contact. In essence, the ground water erodes the underlying shales and undercuts the sandstone. Large sandstone blocks fail in a toppling mode and cause safety and cleanup problems. Use of rock reinforcement including chain link fencing, roof bolt mats, rock bolts, rock anchors, and shotcrete offers a possible solution to the problem.

Jan 1, 1985

By B. K. Parekh

Silicon carbide (SiC) can be produced by heating rice hulls to elevated temperature, The product is a mixture of beta crystalline silicon carbide whiskers, particles and unreacted carbon. The silicon carbide whiskers are 0.5 micron in diameter and 50 to 100 micron in length and have been found useful for reinforcing metal and other composite materials. The carbon can be removed by combustion, but this makes separation of the silicon carbide whiskers from particles difficult and a high grade whisker product cannot be made. However, a clean high grade Sic whiskers product can be obtained from the pre-burned converted rice hulls using physical separation processes that are based on differential wetting. Recovery of Sic whiskers from the pre-burned material was found to exceed 90 percent. A modification of the froth flotation process developed during this study is now being used commercially to separate silicon carbide whiskers and particles and to recover a high grade whisker product.

Jan 1, 1985

By I. Iwasaki, K. Adam

The oxidation behaviors of simple and complex chloride systems of copper, iron, nickel, cobalt, and manganese in a stream of oxygen were investigated using a thermogravimetric analysis technique. Their reaction mechanisms are discussed from stoichiometry, thermodynamic feasibility, and X-ray diffraction analysis of reaction products. The results are used to determine an optimum operating temperature of the selective oxidation step in a chlorination process for producing nonferrous metal chlorides, elemental sulfur, and ferric oxide from copper-nickel sulfide flotation concentrates.

Jan 1, 1985

By W. Weissberger, Frankiewicz, L. Pommier

Introduction In the U.S., about one-quarter of the fuel oil and natural gas consumption is associated with power production in utility and industrial boilers and process heat needs in industrial furnaces. Coal has been an attractive candidate for replacing these premium fuels because of its low cost, but there are penalties associated with the solid fuel form. In many cases pulverized coal in unacceptable as a premium fuel replacement because of the extensive cost of retrofitting an existing boiler designed to burn oil or gas. In the cases of synthetic fuels from coal, research and development still have a long way to go and costs are very high. Another option, which appears very attractive, is the use of solid coal in a liquid fuel form - coal slurry fuels. Occidental Research Corp. has been developing coal slurry fuels in conjunction with Island Creek Coal (ICC), a wholly-owned subsidiary. Both coal-oil mixtures and coalwater mixtures are under development. ICC is a large eastern coal producer, engaged in the production and marketing of bituminous coal, both utility steam and high quality metallurgical coals. There are a number of incentives for potential users of coal slurry fuels and in particular for coal-water mixtures (CWMs). First, CWM represents an assured supply of fuel at a price predictable into future years. Second, CWM is available in the near term; there are no substantial advances in technology needed to provide coal slurry fuels commercially. Third, there is minimal new equipment required to accommodate CWM in the end-user's facility. Fourth, CWM is nearly as convenient to handle, store, and combust as is fuel oil. Several variants of CWM technology could be developed for different end-users in the future. One concept is to formulate slurry at the mine mouth in association with an integrated beneficiation process. This slurry fuel may be delivered to the end-user by any number of known conveyances such as barge, tank truck, and rail. Slurry fuel would then be stored on-site and used on demand in utility boilers, industrial boilers, and potentially for process heat needs or residential and commercial heating. An alternative approach is to formulate a low viscosity pre-slurry at the mine mouth and to pipeline it for a considerable distance, finishing up slurry formulation near the end-user's plant. Finally, at the other extreme of manufacturing alternatives, washed coal would be shipped to a CWM manufacturing plant just outside the end-user's gate. Depending on fuel specifications and locations of the mine and end-user facility, any of these alternatives may make economic sense. They are all achievable in the near term using existing technology or variants thereof. The Coal-Water Mixture CWMs contain a nominal 70 wt. % coal ground somewhat finer than the standard pulverized ("utility grind") coal grind suspended in water; a complex chemical additive system gives the desired CWM properties, making the suspension pumpable and preventing sedimentation and hardening over time. Figure 1 shows the difference between a sample of pulverized coal containing 30 wt. % moisture and a CWM of identical coal/water ratio. The coal sample behaves like sticky coal, while the CWM flows readily. The combustion energy of a CWM is 96-97% of that associated with the coal present, due to the penalty for vaporizing water in the CWM. Potentially any coal can be incorporated in the CWM, depending on the combustion performance required and the allowable cost. CWMs are usually formulated using high quality steam coals containing around 6% ash, 34% volatile matter, 0.8% sulfur, 1500°C (2730°F) initial deformation temperatures, and energy content of 25 GJ/t (21.5 million Btu per st). Additional beneficiation to the 3% ash level can be accomplished in an integrated process. There are a number of minimum requirements which a satisfactory CWM must meet: pumpability, stability, combustibility, and affordability. In addition, a CWM should be: resistant to extended shear, generally applicable to a wide variety of coals, forgiving/flexible, and compatible with the least expensive processing. It was found that a complex chemical additive package and control of particle size distribution are necessary to achieve these attributes simultaneously, while maximizing coal content in the slurry fuel. Formulation of Coal-Water Mixtures A major consideration in the manufacture, transportation, and utilization of a slurry fuel is its pumpability, or effective viscosity. Most CWM formulations are nonNewtonian, i.e., viscosity depends on the rate and/or duration of shear applied. Viscosities reported in this paper were obtained using a Brookfield viscometer fitted with a T-spindel and rotated at 30 rev/min, thus they are apparent viscosities measured at a shear rate of approximately 10 sec-1. The instrument does reproducibly generate a shear field if spindle size and rotation rate are held fixed. By observing the apparent viscosities of several slurries at fixed conditions it is possible to obtain a relative measure of their viscosities for comparison purposes. A true shear stress-shear rate relationship at the shear rates at which the CWM will be subjected in industry may be obtained using the Haake type and a capillary viscometer. These viscometers are used for specific applications. However, for comparison purposes, apparent viscosities are reported.

Jan 1, 1985

The original leases making up today's North Kalgurli Mines were developed by independent companies on an ad hoc basis. The major disadvantage which now arises is the varying levels of mining horizons. From Kalgurli Shaft in the south, through main Shaft and to Croesus Shaft in the north, there are three separate sets of levels. work has commenced to link the three areas with either common levels or new development which will facilitate common access and haulage via Main Shaft.Generally, the workings at North Kalgurli were established at the turn of the century. The introduction of techniques and equipment to bring the mine into the 20th Century has been difficult and slow. The general size of openings, a widespread workforce and complex ore distribution often make innovation and change more than a challenge.

Jan 1, 1985

Gallery Stoping, an old but little used stoping method, has been reintroduced to workings at the Paringa South Shaft.Orebodies which are amenable to such a method are being mined in this manner with considerable advantages and savings.The advantages of gallery stoping as used at Paringa, compared with traditional mining methods are:1. constant cash flow2. low set up cost and initial stope development3. nearly 1005 recovery 4. no access problems 5. quicker tonnagesAs compared with highly mechanized mining methods, it has been found to be slower but requires less development, gives improved grade control, has less impact on the surrounding rock and fragmentation is much better, reducing secondary blasting costs.This paper examines the mining method and illustrates the advantages of gallery stoping at Paringa.

Jan 1, 1985

By David Lynn Passmore

The reliability of mine operators reports to the Health and Safety Analysis Center of the U.S. Deparetment of Labor of days lost from work by injured miners is .267.

Jan 1, 1985

By C. L. Lin

Linear grade distribution measurements determined from polished sections of randomly oriented particles can be used to estimate the extent of liberation for multiphase particles of specified size. Two techniques, optical image analysis and microprobe analysis, have been used to establish linear grade distributions. Both techniques were evaluated and compared using variable grade samples of products from a copper porphyry flotation circuit. The statistical significance of the linear grade distribution data obtained by optical image analysis was assessed. The average chalcopyrite content of these samples, as determined by chemical assay was used to estimate the variance of the experimental procedure. For a given accuracy, the number of particles required to obtain an unbiased representation of the sample was determined.

Jan 1, 1985

By A. J. Kroha, N. Wesis

Most copper mines produce both ore and low-grade "leach" rock or acid waters that contain recoverable copper. The sulfide ores pre¬dominate, and a portion that is too low grade for milling to produce concentrates for smelting, but has to be mined and trucked away anyhow, may be leached successfully with acid in dumps. Most of this leach material consists of sulfides and silicates or carbonates, and if the gangue is such that it consumes a high quantity of acid, this factor may rule out a leach operation. There are also valuable deposits that contain mostly acid-soluble copper, or occasional sulfide ores from which a sulfide concentrate can be roasted and acid-leached to produce a copper-bearing solution. Finally, there are milling ores in which the lesser part of the copper is acid-soluble and can be precipitated with iron or synthetic inorganic precipitants that produce metallic copper or copper sulfides that will float with the sulfides. Ordinarily, ores that contain copper associated with the sulfur ion, such as in the minerals chalcopyrite, chalcocite, bornite (and others), are milled to produce a 25-30% Cu concentrate for smelting, while a lesser amount of acid-soluble copper may be converted from solution to cement copper on iron scrap. A fast-growing percentage of such copper, however, is removed from solution with exchange resins or organic compounds in organic carriers such as kerosene (solvent extraction), then eluted with strong acid and subjected to electrolytic precipitation either in marketable form or as anodes that can be refined further. From the point of view of conventional copper smelting, copper flotation concentrates and cement copper are of chief interest in this chapter. Table I is a condensed open schedule for concentrates that generally run between 25 and 35% copper, and much less frequently as low as 12-15% or as high as 65-75% copper, the former being due to intimate relationship with pyrite (like the former United Verde Extension), and the latter representing such ores as the Bolivian Coro¬coro ore in which the copper is in the form of chalcocite in sandstone. These extremes are no longer common. When they occur, a special purchase schedule has to be negotiated. Included in Table 1, copper precipitates (cement copper) generally run from 70-85%a copper, and the same basic purchase schedule is used as with flotation concentrates. Sulfide Flotation Concentrates The sulfide copper concentrate produced in the mill as a flotation froth, with water then added for transportation of the heavy mineral particles from the flotation cells to thickeners, may run 60-80% water by weight, and the removal of water down to 25-50% by weight by means of thickeners, followed by further dewatering by continuous vacuum filters to 7-18% moisture by weight (depending on size of solids by screen analysis and also by content of clay) is a critical operation. Mill operators would like to produce a filter cake with 7-9% moisture content, but even with the help of steam on the filter this desirable condition is seldom realized when the concentrate is as fine as 80% -325 mesh. More commonly, the final concentrate is reground in pro- to produce best copper recovery and grade of concentrate (or molybdenite separation). In those cases, increasingly frequent, the filter product may not be a cake at all, but a mud that is hard to handle-even requiring a thermal dryer. Greater difficulty of form¬ing a manageable cake often comes from the copper-molybdenum separation by flotation, because the alkaline sulfides and hydrosulfides, or cyanide, or other similar reagents used for the separation, may leave the now relatively molybdenite-free copper concentrate even more difficult to filter. Handling a wet filter cake is difficult enough when its destination is only a short distance away-a matter of yards rather than miles. In those cases the filter cake may be thermally dried near the point of production, using rotary or multiple hearth, or fluidized-bed dryers. Alternatively, the concentrate may be pumped or carried in slurry form to the smelter and filtered there, or it may be spray-dried and compacted. For transportation to a smelter just a few miles to a few thousand miles away by ship or railroad other factors may be important, such as: in shipping by sea, avoidance of spontaneous combustion; in shipping by rail, losses by leakage if too wet or by wind and sun if too dry. It is the responsibility of the millman-usually the mill superinten¬dent-to make sure that his concentrates are in satisfactory condition when they leave the mill so that they meet these requirements: 1) They must have been accurately sampled and dry-weighed, the latter meaning that a moisture determination and gross weight must have been taken. 2) They must be dried sufficiently when necessary to prepare them for safe transportation. 3) They must arrive at the smelter with reasonable likelihood that they can be check-weighed and sampled fairly and equitably,

Jan 1, 1985

By Charles A. Nork

Licensing of surface blasters as a regulatory requirement has been in effect in Pennsylvania for more than twenty years. Training of license applicants was begun in 1971 on a random basis. Formalization of training of applicants for surface blaster licenses was begun in 1980. Training is conducted on a regularly scheduled basis in specific locations throughout Pennsylvania. Training is conducted by Charles A. Nork with the assistance of one of five Department explosives field inspectors.

Jan 1, 1985

By A. L. Mular

Introduction The cost of grinding may be divided into a fixed equipment cost portion and an operating cost portion. The fixed equipment cost is the amount of money required to purchase the grinding mill itself. This amount may be kept separate from the installation cost which, often, is taken as a percentage of the fixed equipment cost. A good average is 43%; therefore a $100,000 ball mill would cost approxi¬mately $143,000 installed. Operating costs are variable and paid per unit time for power, liner replacement, grinding charge replenishment, labor, and mainte¬nance including major parts replacement. Such costs depend on the geographical location and will vary with the cost of labor. Operating costs are difficult to estimate. Most mining companies consider such information proprietary. This section provides a means to obtain an estimate of the equip¬ment cost of a grinding mill with a rated accuracy of±l0%. Compara¬ble methods to determine operating costs are not available. Such costs are presented in table form with suggestions for updating them. Esti¬mating the cost of liners and steel replacement is possible as described in a later section. Capital Costs of Grinding Mills Grinding mill costs have been related,124 by means of equations and corresponding graphs, to the nominal horsepower requirements of mills. All graphs are based on a Marshall and Stevens cost index for Mining and Milling of 300, where a cost index is a ratio used to estimate current prices of equipment from obsolete prices. Where (Cost Index Now) Cost Now = (Cost Then) (Cost Index Then) (I) For example, the cost of an autogenous mill that requires a 3000¬hp motor is estimated as $500,000 (Fig. 95). This was the estimated price when the M&S (Marshall and Stevens, Mining and Milling) cost index had a value of 300. If the index is currently at 400, then the estimated current price of the autogenous mill is: Cost Now = ($500,000) (400) (300) = $667,000 Most professional journals will publish current values of cost indices at least once a month. In subsequent pages, each graph of grinding mill cost vs. mill horsepower has an associated graph of capacity vs. horsepower along with a table that relates mill size to horsepower. The associated graph and the table must be used with caution, since both apply to an average ore under average conditions. Example calculations showing how graphs should be used are shown in the following. It should be stressed that the cost data incorporated herein ex¬cludes the cost of pumps, motors, lines, and classifiers, unless other¬wise stated. Costs are not direct quotes; they are the averages of quotes supplied by manufacturers and analyzed by means of regression analysis. Example 1. Calculate the power required to grind 100 tons of dry solids per hour (in a closed circuit with a classifier) when the

Jan 1, 1985

By G. V. Sullivan, B. E. Davis, T. O. Llewellyn

The Bureau of Mines, US Department of the Interior, conducted studies to recover phosphate from four dolomitic southern Florida deposits not currently mined. The samples contained 5.8-10.2% P205 and 1.7-4.8% MgO. Chemical analyses of sized fractions indicated that no pebble concentrate, +1.2 mm (+14 mesh) could be produced from these samples. A scheme was developed that consisted of desliming, grinding, scrubbing, and redesliming the ore to produce a flotation feed. Conventional one-step fatty acid fuel oil batch flotation tests on the samples resulted in concentrates containing 29.0-31.3% P205 at flotation recoveries ranging from 72.4 to 96.1% of the flotation feed. The MgO content of the concentrates ranged from 0.47 to 1.36%. Continuous flotation of the Manatee No. 2 sample indicated that commercial-grade phosphate concentrates analyzing 31.3% P205 and 0.55% MgO could be produced with a flotation recovery of 89.9%. The MgO content of concentrates obtained from the Manatee No. 1 and Hardee samples was 1.20 -1.36 %, respectively. Leaching with aqueous SO2 reduced the MgO content of these phosphate concentrates to 0.5 and 0.8%, but with a concomitant loss in phosphate recovery of about 10%.

Jan 1, 1985

By G. B. Lundstrom

SYNOPSIS For some time, Vaal Reefs South has operated blindhole borers successfully as an integral part of the mining cycle of operations. To date, over 13 000 m of mechanized boxholing has been completed, and from time to time impressive performances have been achieved. The paper describes the machines, the modus operandi, and the structure of the boring organization at the Mine, with some of the results recorded. The Mine recently took delivery of two new machines with interesting design features. The introduction and early experiences with these two machines are also described.

Jan 1, 1985

By M. J. Taylor

A common method of classifying non-water- impounding mine waste structures is necessary to facilitate professional comnunications within the industry and between representatives of industry and government. Camnon classifications can a1 so he1 p quickly create and maintain a mental picture of the structures being discussed. Three categories for classification presented in this paper include: (1) configuration and location with respect to existing topography. (2) material characteristics of waste, and (3) environmental sensitivity or characteristic; of the disposal area. Configuration and location classification involve terminologies typically utilized for waste dumps over the past several years. General classification categories for non-impounding structures include: (a) valley-fill, (b) cross-valley, (c) side-hill, (d) ridge, and (e) heaped. Material characteristics can govern the location, engineering details, and environmental sensitivity of the waste structure. Material classification includes: (a) hazardous, (b) non-hazardous, (c) inert, (d) fugitive dust susceptible and (e) leachate susceptible. A1 though many characteristics are potentially applicable, this paper will focus on the categories mentioned above. Environmental sensitivity of a waste structure encompasses many objective and subjective evaluations plus many multi-disciplinary considerations. Environmental sensitivity incorporates numerous factors in addition to those highlighted in configuration and location, and material characteristic classifications. A few examples of how this categorization can govern the engineering design and location of waste dumps are provided in this paper. In addition to the verbal descriptions of classifications, a simplified version of the combined classifications with an alphanumeric format and its methodology is discussed in this paper. Surface water controls often govern the classification or vice versa. In order to maintain a non- impounding status, water must be prevented from collecting behind the structures. Depending on material characteristics, water inflow to or over the site can cause significant environmental or human consequences. Finally, the access of water to the waste dumps can play a large role in the degree of environmental sensitivity to the site. This paper provides a framework in which to discuss the other aspects of this conference and also provides insights as to how the framework is based on actual practice.

Jan 1, 1985